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By William M. Gleason

The war in Vietnam stretched into another decade, an American President was at the brink of impeachment, the earth was given its own day, twin tower monuments of American ingenuity were completed in New York and would stand high above the Manhattan skyline for 31 years. Inflation, oil shortages, low commodity prices and America’s 200th birthday were part of the decade — it was the 1970s. For the mining industry, the demands of the decade would be high and the challenges would be significant. Because of the many challenges that lined up in front of it, the mining industry was forced to work smarter and more efficiently. The industry also expanded its borders in search of new resources and struggled along with the rest of the world through a fuel crisis, inflation and a recession.

Jan 1, 2007

By R. W. Chappell, M. W. Hills, R. L. Olsen

A stochastic approach was developed and applied to the Butte, Montana Hillside abandoned mining site for modeling stormwater runoff and subsequent receiving stream loadings. This approach enabled capture and quantification of the uncertainty associated with stormwater quality data and allowed the prediction of copper and zinc concentrations in the receiving stream caused by runoff from the Butte Hillside during storm events. Runoff flows and stormwater concentrations were input to the model as probability distribution functions. Monte Carlo simulation allowed prediction of expected instream concentrations in the form of cumulative distributions functions. The modeling results allowed evaluation of current best management practices (BMPs), targeting of future BMPs, and identification of key data gaps to fill during future sampling activities. This modeling approach is considered extendable to other active and inactive mining sites. Output from a model of this type is easily understood and can be used to communicate effective stormwater management to site regulators. Furthermore, application of this modeling approach will facilitate the development and implementation of a site-specific stormwater management plan and will optimize efforts to achieve standards for compliance with stormwater discharge permits.

Jan 1, 2002

By Lok W. Home

HISTORICAL DEVELOPMENT At the beginning of the 1970 more than 90% of pro¬duction drilling in open-pit mining was obtained with the use of rotary drills. At the same time, the use of rotary drills underground was virtually nonexistent. Several progressive mines began questioning whether efficient rotary drilling could not be adapted to under¬ground applications. In the late 60s and early 70s at¬tempts to use rotary drilling were made by Inco, Ltd. and Meramac Mining, utilizing drill units designed as a combination of small diameter raise drills and rotary drills. These initial attempts were unsuccessful primarily due to the immobility of the equipment and low degree of dependability. At the same time, experiments con¬ducted by Inco and Meramac using large diameter in¬the-hole percussion drills produced immediate encourag¬ing results. The first serious attempt at utilizing rotary drilling underground was made by Texasgulf in 1971 at Kidd Creek mine in Canada. An experimental stope 15 m wide, 36 m long, and 30 m deep was drilled using 200-mm diam rotary drilled holes. The machine used was a Robbins' Model 11 D. The Model 11D is a rotary drill designed to drill drain and utility holes under¬ground. The results of this test revealed that large diameter blasthole stoping was feasible. The indications were that the major areas of concern-fragmentation and pillar damage-could be controlled and a new low-cost stoping method resulted. At this point Texasgulf and Robbins entered into a cost-sharing agreement to develop a rotary drilling unit specifically for this stoping method. Drilling of the trial stope required pouring a concrete pad at each drill hole plus anchoring the derrick to the pad by means of rock¬bolts. Utilization expressed in terms of drilling time to available time was less than 20%. It was obvious that the required unit had to be mobile, setup time had to be minimized, and rod handling had to be improved. The first attempt at an underground production-type rotary drill consisted of mounting the basic 11D on an air-powered Eimco 600 Series crawler. The crawler was equipped with four screw-type leveling jacks operated by an impact wrench; rod rack facilities for 60 m of 1.2-m long, 170-mm diam rods; a rod handling facility consisting of an air-lift swing boom and a pipe loader which automatically positioned the rod under the float¬ing drive box; a turnbuckle anchor system requiring in¬stallation of four rockbolts in the drill drift floor and placing turnbuckles from the rockbolts to the corners of the drill; and trailing power units containing the electrics and hydraulics required to operate the drilling unit. Operation of this unit revealed a vast improvement in drilling time but some problem areas remained. The anchoring system required secondary drilling and proved difficult to adapt to the localized irregular floor surface and the trailing power unit was difficult to maneuver and was causing unnecessary maintenance costs. Addi¬tional improvements resulted in the standard 11MD rotary blasthole drill. DESCRIPTION OF UNIT This standard model consists of the drill and the power unit integrally mounted on an air-powered crawler chassis (Fig. 1). The rod-handling and leveling¬jack systems are similar to those used on the prototype Texasgulf unit and remained relatively the same. The anchoring system has been replaced by the roof jack, initially screw type but now hydraulically actuated. The anchoring system is necessary as the drill exerts more thrust than the weight of the unit. During operation it is found that the two front jacks adequately anchor the machine. The complete unit's operational dimensions are: height, 3.2 m; width, 2.1 m; and length, 4.2 m requiring a 3.5- x 3-m drill drift. The derrick is pivot mounted and raised and lowered by a hydraulic cylinder reducing the transport height to 2.5 m. The derrick is also turnbuckle mounted to allow dip angle drilling from vertical to 0.52 rad (30°) off vertical. The drill derrick can be mounted to allow dip axis to be parallel or perpendicu¬lar to the long axis of the unit. The standard rods are 1.2 m in length and are contained in racks of 10 which can be removed for transportation. Drilling torque is produced by a 56-kW (75-hp) electric motor driving a double vane pump which in turn drives a Robbins high torque gear-type hydraulic mo¬tor. In addition to the hydraulic motor, the machine drive system has a 3.9:1 planetary gear reducer. This results in six speeds with the following minimum and maximum range: Maximum rpm 112 Torque 3725.5 J (380 kg m) Minimum rpm 10 Torque 14 706.0 J (1500 kg m) Drilling thrust is provided by two bootstrap cylin¬ders pressurized by a variable volume pump. This sys¬tem can produce from 0 to 267 231.2 N thrust in either direction. Bailing air (approximately 17 m3/min at 382.2 to 490.3 kPa) is introduced through a rotary swivel pass¬ing directly through the center of the drive system to keep efficiency and losses to a minimum. Hole annulus is also kept to a minimum by using 170-mm rod when drilling a 200-mm hole. A small amount of water is added to the air for dust suppression. The machine has to be supplied with air, water, and electrical power. It does not appear feasible to provide an integral air power supply with the rotary drill such as is done with open-pit rotary drills. Air is required and installed during the primary development of the drill drift. The machine can be adapted to drill hole diameters from 170 to 228 mm for blasthole operations. The drill string consists of the rotary bit, crossover stabilizer, roller stabilizers, and standard rods. The operator controls speed and thrust pressure from an operating console which can be mounted in a conve¬nient position. Mount Isa has made an adaption to the operating console to permit remote collaring of drill holes in an uncleared floor.

Jan 1, 1982

By MOHAMED WAEL, ANASTASIA GOEVA, ABDULAZIZ TAWFEEQI, IHTISHAM TARIQ, S. M. ZAKIR HOSSAIN, NIAZ ALI KHAN, MUHAMMAD FAISAL IRFAN

Carbon dioxide (CO2) mineral carbonation is a carbon capture and storage (CCS) technique that can sequester CO2 permanently in the form of durable carbonate minerals. Extendable permanent storage capacity — onshore and offshore storage reservoirs — and the exothermic nature of the mineralization process make this technique attractive [1]. In the present study, maximum carbonation efficiency for aqueous mineral carbonation was achieved through two steps: optimization of calcium leaching from cement kiln dust followed by the reaction of pure CO2 with the calcium hydroxide precipitates formed by hydroxylation using sodium hydroxide (NaOH). Optimization of calcium leaching is highly important, as it is a rate-limiting reaction step in mineral carbonation and also influences and enhances the carbonation efficiency. To maximize the calcium leaching as well as carbonation efficiency, response surface methodology — with either Box-Behnken design or central composite design — was applied. It is the most well-known statistical, fully randomized and bias-free method for determining optimized conditions using quadratic regression models.

By Erika Moonin, Jim Nickerson, Marcus Jensen

"This paper will explain how the Contractor and Owner avoided a contractual dispute involving characterization, impacts, and final remedy for challenging ground conditions encountered during drill and blast excavation of the starter tunnel required for launch of the tunnel boring machine (TBM). During excavation more than 600 feet underground, the Contractor encountered a large complex fault zone subject to more than 14 bar of water pressure. Successive ground inflows from the fault complex over a six-month period presented significant technical and partnership challenges for the team and ultimately delayed the project by nearly one year. This paper will discuss how the Contractor, Owner, and their respective consultants worked together to address these challenges, make decisions, and ultimately move the project to successful completion. PROJECT BACKGROUND AND NEED Seven states of the western United States of America rely heavily on the Colorado River and its tributaries for their water supply. The primary source of water for the Colorado River is winter snow that falls in the mountains ringing the Colorado River basin, including the Rocky Mountains, the Wind River Mountains, and the Uinta Mountains. Flows in the Colorado River vary significantly from year to year, depending on the amount of snow and rain deposited in the basin. During the 1980s and 1990s, snowfall was relatively abundant and Colorado River flows were high on average. Beginning in the year 2000 and continuing to the present day, snowfall has been well below documented historical averages and Colorado River flows have diminished. In fact, a trend toward diminishing flows is evident during the period of the historical record. This variability in Colorado River flows and the trend to lower flows is represented in the following Figure 1."

Jan 1, 2016

By Richard J. Wilson

The importance of coal as a major source of energy used by the steel industry has been increasing due to the gradual changeover from steel- making by the open hearth process to steelmaking by the basic oxygen process. Since the production of steel by the BOF process increases the demand for hot metal, there is a growing need to ensure the supply of coke required for the production of hot metal. Coal used for cokemaking has become, therefore, an increasingly important raw material for the steel industry in recent years. Our current projections indicate that coal will be the source of 71% of the total energy requirements of Inland Steel Company in 1985. As a result of the energy shortages of the last few years, various conservation programs have been initiated by the steel industry to reduce the consumption rate of fuel. The efforts to conserve energy have been successful to a great extent due to the rapid advances made in the technology of ironmaking that facilitated gradual reductions in the quantity of coke used per ton of hot metal produced in the blast furnaces. The decrease in the consumption rate of coke was made possible by the use of improved raw materials, such as iron ore pellets and fluxed agglomerates, the operation with high blast temperatures, the injection of auxiliary fuels, and the enrichment of the blast with oxygen. It is expected that additional reductions in the fuel consumption rate will be possible in the ironmaking operations of the large blast furnaces that are being equipped with facilities for proper burden distribution and for operation with higher blast temperatures at high top pressure. The purpose of this paper is to show the effects of coke quality on the production of BOF steel. The importance of coke quality will be pointed out by discussing the effects of the physical strength of coke on both the blast furnace productivity and the steelmaking operations of an integrated steel plant, such as Inland Stee1 Company.

Jan 1, 1980

Filtration systems with disposable filter elements (DFEs) are used in the underground coal mining industry to control particulate matter emissions from diesel-powered permissible and nonpermissible coal mining equipment. This study was conducted in underground mine conditions to evaluate three types of high-temperature DFEs used in those filtration systems. The DFEs were evaluated for their effects on the concentrations and size distributions of diesel aerosols and concentrations of nitric oxide (NO) and nitrogen dioxide (NO2). Those effects were compared with the effects of a standard muffler. The experimental work was conducted directly in an underground environment using a unique diesel laboratory developed in an underground experimental mine. After an initial DFE degreening period, the filtration system with all three DFEs was found to be very effective at reducing total mass concentrations of aerosols in the mine air. The effectiveness of DFEs in filtering aerosol mass was found to be a function of the engine operating conditions. The efficiency of the new DFEs significantly increased with accumulation of operating time and buildup of diesel particulate matter in the porous structure of the filter elements. A single laundering process did not exhibit substantial effects on performance of the DFE elements. The effectiveness of DFEs in removing aerosols by number was strongly influenced by engine operating mode. The concentrations of nucleation mode aerosols in the mine air were found to be substantially higher for both DPFs and DFEs when the engine was operated at high-load modes than at low-load modes. Initial heating of certain DFEs resulted in visible white smoke and substantially elevated aerosol number concentrations. The effects of the DFEs on total nitrogen oxides (NOX) concentrations were found to be minor. The NO2 fraction was found to be generally lower for the DFEs than for the muffler. The engine-out NO2 fraction of the total NOX was found to be substantially higher for low-load modes than for high-load modes.

Jan 1, 2011

By M. A. Trevits

The National Institute for Occupational Safety and Health (NIOSH), in partnership with the Mine Safety and Health Administration (MSHA), is conducting research to test, evaluate, improve or modify coal mine fire-fighting strategies and methodologies through large-scale tests. Since wood is the second most abundant fuel available during a coal mine fire, it was decided that series of large-scale wood crib fire tests were needed to measure the products of combustion and to observe the capability of commonly available direct application fire suppression equipment, namely, fire extinguishers, water and gas-enhanced foam. This paper discusses the wood crib fire tests, provides insight into the products of combustion and describes observations made during the application of the fire suppression agents.

Jan 1, 2009

By Tom Ireland, Sumi Eratne

The approach to Risk Management during the procurement of major projects often has a significant outcome on the ultimate successful delivery of a project. The $NZ4.4B Auckland City Rail Link project is the largest and most complex transportation project yet undertaken in New Zealand. Risk management prior and during procurement was undertaken in accordance the ITIG Code of Practice for Risk Management of Tunnel Works (TCoP) and through the use of an Alliance Geotechnical Baseline report. This paper describes the measures taken during the procurement process to achieve certainty with respect to outcomes and minimise project risks whilst also achieving value for money. This included the use of cloud-based delivery of geotechnical data and 3-D geological models, reducing tendering costs and allowing a focus on risk mitigation measures during the tender rather than data management.

Jun 13, 2021

By J. Noll, J. A. Organiscak, A. B. Cecala

"In recent years, significant strides have been made to optimize the design of filtration and pressurization systems used on enclosed cabs of mobile mining equipment to reduce respirable dust and other contaminants and provide the best air quality to the equipment operators. This synergetic effort has involved original equipment manufacturers building mining equipment, manufacturers of heating, ventilation, and air conditioning systems for mobile equipment, filtration and pressurization system manufacturing companies, as well as the National Institute for Occupational Safety and Health, and other government organizations. Considering all the advances made in this area over the last decade, one aspect that NIOSH still needs to assess is the longevity of optimal cab performance with respect to the quality (filtering efficiency) of the filters used in filtration and pressurization systems. In today’s culture, most health and safety professionals automatically believe that HEPA quality filters need to be used to provide the greatest level of protection to workers. Researchers at the Office of Mine Safety and Health Research for NIOSH hypothesized that HEPA quality filters may not be optimal for the mining industry and speculated that MERV 16 rated filters would be more appropriate in most cases. In order to test this hypothesis, NIOSH performed a two-year study comparing both HEPA and MERV 16 quality filters on two pieces of underground limestone mining equipment, being a roof-bolter and face drill machine. Testing showed at the 95 pct. confidence level that there was no statistical difference between the two efficiency filters on both pieces of mining equipment. Since the MERV 16 rated filters were less restrictive and provided greater cab pressure, and since they did not have to be replaced as often as the HEPA quality filters, researchers concluded the MERV 16 filters were the optimal choice for both pieces of mining equipment in this case study comparative analysis. INTRODUCTION When most health and safety professionals today think about filtration efficiencies and their correlation with protecting workers’ health, the normal assumption is the higher the efficiency of a filter, the greater the protection afforded to the workers. The next logical step is to believe that HEPA filters deliver the greatest protection for workers because they provide the highest filtering efficiency. Obviously high-efficiency intake filters are a necessity for an effective cab filtration and pressurization system on mobile mining equipment, but what is the optimal filter efficiency rating for achieving high levels of cab protection factor performance over the service life of filters? To address this question, the National Institute for Occupational Safety and Health (NIOSH) has performed an in-depth laboratory research study over the years simultaneously with numerous field studies to retrofit used enclosed cabs on mobile mining equipment with newer and more effective filtration and pressurization systems."

Jan 1, 2016

By Paul Dean Proctor

The world's population by 2000 A. D. may approximate 6.5 billion and the United States' 304 million. South Asian population increases between 1970 and 2000 A. D. will be triple the projected population of North America. Such anticipated dramatic and probably traumatic increases in world population yield grave concern for the standard of living in the world and the United States. The people of the United States currently consume 32 percent of the world's energy and similarly large percentages of the world's mineral raw materials. With only 6 percent of the present world's population and slightly smaller percentage by 2000 A. D., much expanded world-wide production of energy and raw materials will be required to maintain even a standstill living standard. A minimal 2 percent increase per year in the standard of living in the United States and a 4 percent increase elsewhere in the world, beyond the demands of the estimated population increases, suggest the magnitude of the increased need for energy-mineral raw materials and the possible crises the world will face in these areas by 2000 A. D. Mineral engineers and scientists with broad educational backgrounds will be required in greatly increased numbers to help alleviate, and possibly avert, expected energy-mineral raw materials shortages of the next millenium. Explorationists, mineral engineers, commodity economists, environmentalists, city, county, state, and federal mineral materials planners and coordinators, and metallurgists will be among those most needed. AIME membership in the second century of the Institute's history will contribute directly to this personnel pool.

Jan 1, 1971

By R. M. Rowe, A. K. Bhasin, D. Malhotra

Introduction Most primary molybdenite producers use a selective flotation process in which molybdenite is floated from the ore while other sulfide and nonsulfide gangue minerals are depressed in the same step. The process was developed at the Climax mill. Reagents used in the rougher flotation stage include vapor oil (collector), syntex (surfactant), pine oil (frother), and occasionally sodium cyanide and Nokes reagent (depressant). Rougher concentrate is upgraded by cleaner flotation to produce market-grade molybdenite concentrates. Reagents used in the cleaner flotation stage are Dowfroth 250, vapor oil, sodium cyanide, and Nokes reagent. The development of the flotation reagent scheme at Climax and the role of the reagents have been well described by Born et al. (1976) and Malhotra et al. (1976, 1977a, 1977b, 1977c, 1980). The hydrocarbon collector, vapor oil, accounts for approximately 50% of the total molybdenite flotation reagent cost. Because vapor oil sources are limited, and the cost of vapor oil has increased significantly over the last decade, several studies have been done to evaluate alternate oil-based collectors and/or blends of collectors that would produce similar metallurgical results at a reduced cost, or increase the overall molybdenite recovery at the same cost (Smit and Bhasin, 1983; Hahn, 1969; Smith and Bhappu, 1971). The approach taken in the present study was to evaluate alternate nonoil-based collectors and collector systems that are less expensive than vapor oil. A series of potential nonoil-based collector classes were identified and evaluated in the laboratory against the current collector system. The metallurgical performance of these collector classes was related to their chemical structures. The results of this study are discussed. Collector classes A survey of sulfide mineral collectors was performed with the objective of identifying and selecting potential collectors for molybdenite flotation. The available reagents were categorized into 10 general classes. Table 1 summarizes these collector classes, their chemical formulas, the manufacturers, and selected commercial collectors in each class that were evaluated in the present study. Experimental procedure A Climax ore sample, assaying 0.3% MoS2, 2.2% FeS2, 0.007% Cu, 0.003% Pb, and 0.03 oz per ton Ag, was used for this study. Standard laboratory rougher flotation test procedure involved grinding a 2-kg ore charge in a rod mill at 60% solids. Reagents were added to the rod mill including a collector, syntex (0.005 kg/t), pine oil (0.03 kg/t), lime (to adjust pH to 8), and sodium silicate (0.15 kg/t). The ground pulp was transferred to a 1000-g cell, and flotation was performed at 35% solids and 1800 rpm using a Denver D-1 flotation machine. Concentrate was collected for a flotation time of six minutes. The general approach used in the present study was to substitute the nonoil-based collectors at two to three different dosage levels (0.01 to 0.15 kg/t) for vapor oil in the bench-scale, open-circuit rougher flotation tests

Jan 1, 1987

By Peter J. Tarkoy

A great deal of attention has been directed toward the complexities and inadequacies of exploration and pre-bid geotechnical information for heavy construction projects, differing site conditions, and the adversary relationships which generally develop as a result. The adversary position has been maintained at all cost. The solution is simple, if not easy. The solution requires strict and equal adherence to given principles by all parties involved, throughout the life of the project, from conception to completion. It is necessary to review and redefine some basic concepts, without value judgements, from a context of getting the job done. In a recreated context, the role of truth, judgement, responsibility, and integrity, in exploration, engineering design, construction estimating, and project construction, will be examined. It is suggested that anticipated geotechnical conditions be defined unconditionally. The alternative choice is to give the bidder the opportunity of making the interpretive judgements based on geotechnical data. Similarly, a unilateral method of establishing encountered conditions that are agreeable to both owner and contractor, must be outlined prior to contracting. As anticipated and encountered conditions are firmly and accurately established, significant differences can evaluated based upon limitations specified in the contract documents. Impact of differences on project costs can be fixed by pre-defined methods.

Jan 1, 1983

By R. K. Singhal, H. Sahay

Introduction In terms of the world's recoverable coal reserves, Canada ranks 10th, providing 0.6% of the world's total. The US is first with 27.8% and the Soviet Union follows with 24.1%. Figure 1 shows the principal coal mines in Canada. Several provinces contain substantial coal deposits, but the largest deposits are found in the western provinces of Alberta, British Columbia, and Saskatchewan. Conservatively, 90% of the lignitic, 80% of high volatile bituminous, and nearly 100% of subbituminous, and low and medium volatile bituminous coals are located in these three provinces. During the past decade, Canadian coal production of all types increased from 18.8 Mt (2 million st) in 1972, to 25 Mt (28 million st) in 1975, to 61 Mt (66 million st) in 1984. About 38 Mt (43 million st) of this output was classified at thermal and 23 Mt (25 million st) as coking and metallurgical coal (Table 1). The three western provinces currently produce 57 Mt (62 million st) of coal - 95% of Canada's total. The remaining 3.6 Mt (4 million st) (5%) is produced in Nova Scotia and New Brunswick on the east coast. The majority of western coal (98%) is produced by surface mining methods. On the east coast, though, mining is predominantly underground and only 15% is produced on the surface. In 1985, Canada exported close to 28 Mt (30 million st) of metallurgical coal and thermal coal. During the same period, 15 Mt (16 million st) of coal were imported, mostly from western US for use in Ontario and Quebec. Almost 60% of the imported coal was used for thermal applications. The remaining 40% was metallurgical coal for the steel industry. It is generally cheaper to import coal into central and eastern Canada from the eastern US to augment local supplies rather than from western Canada. This is mainly because of high transportation costs. Mining methods The suitability of mining methods in a particular area depends largely on the geological setting. Geology, therefore, simplifies or adds complexities to mining operations, directly affecting the economics. It also influences the flexibility required from the min-

Jan 9, 1986

By Louis H. Meece

The romance of the River lives on in the tales of Mark Twain; and many Americans still think of the Mississippi River in terms of the steamboats loaded with passengers, gamblers and cotton. Yet for many years, coal has been a major cargo on the American waterways. Currently, about 100 million tons a year of coal is handled in some part of its journey from mine to destination by barge. The bulk of this traffic has been on the Ohio River System and the Upper Mississippi River System in years past. Now we have a new entry. Over the past 10 years various movements of coal have developed down the Mississippi River to power plants at Memphis and along the Gulf Coast, and a small amount of metallurgical coal has moved down the Mississippi River for export to Japan since 1970. But in 1980 a brand new player came on the scene. 1980 was a momentous year for the American coal business. Historically, we have exported about 50 million tons a year of metallurgical coal to the world's steel mills. Just before and after World War I, we also had a fairly healthy export trade in steam coal; and again in the late forties, we supplied large amounts of coal to Europe until domestic production could recover from the ravages of World War II. But since the early fifties, the name of the export game has been metallurgical coal. Starting in the mid seventies, something began that was almost totally missed by the American coal industry; and that was the growth in the use of coal first in Europe and now in the Far East as a steam product and as an alternative to petroleum products. As these world coal markets began increasing, the U.S. at first was thought to be noncompetitive because ocean rates were still very low. For example, as recently as 1977, Australian coal could be moved all the way from Australia to the Gulf Coast for $12.50 per metric ton. Coal exporting countries were Poland, Australia and South Africa, except for the continuing exports of U.S. high quality metallurgical coal. But then came the second oil shock, and all of the sudden that $12.50 rate from Australia to the U.S. Gulf became $25.00. And then suddenly and almost unrecognized by the American industry, American coal was in fact competitive at overseas destinations because of the drastic change in trade patterns brought on by the doubling and tripling and quadrupling of bunker fuel prices for ships.

Jan 1, 1982

By H. C. Rolseth

The first step in the quarrying of granite, following removal of any overburden material, involves the freeing of large blocks from the formation. For many years the use of explosives was the accepted method of obtaining these blocks. Even today there are operations where this is considered to be the most economical method even though it does result in a high percentage of damaged stone. In more recent times the preferred procedure at most quarries has been to produce a continuous vertical cut or slot completely around the block to be separated. Originally, this was done by drilling a series of closely spaced holes around the block and breaking the web between the holes with a chisel-like tool known as a broaching bar. The machine used in this mechanical method of producing the relief slot is known as the channel bar. The modern day approach to the production of continuous channels was introduced to the industry around 1955 by the Linde Division of Union Carbide Corporation. The method employs a small rocket - jet burner producing a high temperature, high velocity exhaust flame which when directed against the granite causes it to disintegrate. This method, which has come to be known as "Jet Channeling", is now used by practically every major granite producer in the United States and Canada.

Jan 1, 1968

By S. L. Barua

An industrial engineering and computer simulation study was conducted on reclamation operations to evaluate the possibility of integrating mining and reclamation operations, and to evaluate the factors which influence the production potential of scrapers. The study was done in a southern Illinois strip coal mine producing approximately 2.27 M metric tons (2.5 million short tons) of clean coal annually. The main outcomes of this study are (a) it is possible to reduce reclamation cost about 20-25 percent by integrating reclamation and mining operations, (b) haul distance has the most significant effect on scraper production, (c) push-pull is the best mode of operation for wheel scrapers for topsoil/subsoil removal, and (d) historical data obtained from the mine for one year show that average dozer productivity in rock grading is about 60 percent lower than the expected production; similar data for scraper is about 25 percent less.

Jan 1, 1981

By Lance Rasband, Richard Hanke, John Starcevich

"Mining on Seattle’s SR99 tunnel began in July of 2013. Initial mining progressed slowly through a 4.5 m (14.9-ft) thick fiber-reinforced concrete headwall before immediately entering the first of three planned safe havens. Once through the headwall and safe haven the machine would move through roughly 122 lineal m (400 ft) of jet grout improved soil at a rate of slightly more than 1.6 m/d (5.3 ftpd). It was anticipated that once mining has passed through the improved area, “Bertha” – now the world’s second largest tunnel boring machine (TBM), would be moving at a rate of almost 11 m/d (36 ftpd). By the end of October 2013, Bertha had mined roughly 131 m (430 ft), averaging 1.4 m/d (4.7 ftpd), and had begun mining within native soils. By the start of December 2013, tunneling had reached the 244 m (800 ft) mark and Bertha was averaging 3.6 m/d (11.9 ftpd), achieving as much as 12.8 m (42 ft) of progress on given days. Unfortunately, on Dec. 6th, 2013, Bertha began to stumble and all progress quickly came to a halt.Initial reports stated an obstruction had stopped Bertha in her tracks. The preliminary investigation involved installation of 10 deep dewatering wells to reduce the hydrostatic pressure within the ground around the TBM, to allow the contractor’s personnel to inspect the cutter head from within and also to investigate for the presence of obstructions at the face of the cutter head. During inspection, the contractor’s crews found fragments of steel pipe entangled in the cutter head. The steel was apparently from an abandoned 254 mm (0.9 in.) diameter monitoring well casing, previously installed by the Washington State Department of Transportation (WSDOT) to monitor groundwater conditions for the design and planning of this tunnel project. After the steel was removed from the cutter head and cleared from Bertha’s path, she attempted to move forward but was only able to generate a slow crawl."

Jun 1, 2016

By G. Bulut

The recovery of copper and cobalt from ancient copper slags from the Küre region of Turkey was investigated. A fayalitic-type of ancient copper slag containing 1.24% Cu, 0.53% Co and 53.16% Fe were subjected to an experimental study. Two different routes were followed to recover the metal values. The first route was roasting the slag with pyrite followed by leaching. The second route was flotation of the slags, roasting of the flotation tailings with pyrite and leaching. The second route was found to be suitable for the treatment of the copper slag studied. In the flotation step, a copper concentrate containing approximately 11% Cu was produced with a 77% recovery while 93% of the cobalt stayed in the tailings. In the roasting experiments, the effects of roasting time, pyrite: slag ratio and roasting temperature on the dissolution efficiencies of Cu and Co were investigated, and the optimum conditions were determined. In the tests, 87% of the cobalt dissolution was observed after one hour of roasting at 500°C with a pyrite: slag ratio of 3:1. The copper dissolution in the tests was 31%. Based on the results of this study, a process flowsheet is proposed for the evaluation of these slags.

Jan 1, 2007

By J. W. Powell

Several years prior to 1962, it became obvious that in order to develop ore reserves below the Storke Level, it would be necessary to devise a method to facilitate the ore being lifted three hundred feet to the surface or six hundred feet to the mill. The final plans called for an underground crusher complex with two stages of crushing and a conveyor belt gallery for transport of the ore six hundred feet above, to the third crushing stage. Access into the area was gained in 1962, by sinking Five Shaft approximately one quarter mile southwest of the proposed complex site, and drifting into the area. Collar elevation of Five Shaft is three hundred feet above the 600 Level Crusher Complex. All muck was loaded into ten-ton granby cars, hoisted to the surface and dumped. (See Figure 1 for the relationship between Five Shaft and the Crusher Area) Drifting into the Crusher Area contunued until 1965 at which time the Crusher Room Cutout was started. The Crusher Access Drift (CAD on Figure 2) war driven and mining pm- ceeded according to the description on drawings marked Figure 3a and 3b. After the arch portion of the Cutout was excavated and concreted, removal of the center "pillar" commenced. This muck was removed with an Eimco 123C crawler-mounted front-end loader. (See Figure 4) A total of 576,000 cubic feet of rock was removed from the Crusher Cutout. All sup- port was afforded by grouted roofbolts on a specified pattern, as shown in Figure 3c. Crane support columns were put into place and held by braces welded to roofbolt heads. Sidewalls and endwalls were formed and concreted. (See Figure 5 for the finished product)

Jan 1, 1970